Atmospheric mineral leaching process

ABSTRACT

A method is provided for processing a sulphide mineral composition which at least partly comprises an iron-containing mineral. The method includes the steps of: (a) milling said composition to a particle size P80 of 20 microns or less; (b) leaching said composition with a solution comprising sulphuric acid and ferric ions at ambient pressure while sparging with an oxygen-containing gas in an open tank reactor at a temperature of up to about the boiling point of the solution, whereby at least some of the acid and at least some of the ferric ions are obtained from dissolution of the iron-containing mineral, and ferrous ions generated by the leaching reaction are substantially re-oxidised to ferric ions in the leaching solution; (c) precipitating excess iron and separating said iron together with any solid materials from the leaching solution; (d) extracting desired metal ions from the leaching solution by solvent extraction with an organic solvent, such that the raffinate comprises sulphuric acid and ferric ions; (e) returning the raffinate to the leaching tank and blending with further milled composition; and (f) separating the metals from the organic phase obtained in step (c) by stripping with electrolyte and electrowinning.

FIELD OF THE INVENTION

This invention relates to a method of enabling a sulphide mineralcomposition to be leached at atmospheric pressure instead of aboveatmospheric pressure which has been hitherto required in order toachieve acceptable rates of leaching.

BACKGROUND ART

Sulphide minerals such as copper, nickel, zinc, gold and the like arerecovered from their ores by a number of well known processes. One suchprocess uses the relative solubility of the mineral in solution to allowthe mineral to be leached from the ore. Conventional leaching processesrequire expensive equipment and a high level of technical expertise tomaintain and use the equipment. Thus, it is not uncommon for anoxidative hydrometallurgy leaching plant to be located some distanceaway from the ore body and even in another country. This in turnsignificantly increases transportation costs, and it should be realisedthat transportation of ore or only partially enriched ore containingperhaps only a few percent of the desired mineral is extremely wastefuland undesirable, but in the absence of being able to recover the metalof value from the minerals on-site, there is little real alternative.

The processing methods of oxidative hydrometallurgy are commonly used inmany different applications. Due to the refractory nature of many of themineral species treated in such processes, these applications generallyrequire leaching conditions of high temperature and pressure and requiresubstantial supplies of oxygen. For example, base metals such as copper,nickel and zinc can be recovered by hydrometallurgical processes whichusually embody pretreatment, oxidative pressure leaching, solid/liquidseparation, solution purification, metal precipitation or solventextraction and electrowinning.

According to conventional technology, oxidative leaching processesusually require very aggressive conditions in order to achieveacceptable rates of oxidation and/or final recoveries of metal. Underthese conditions, which often mean temperatures in excess of 150° C. oralternatively temperatures in the range 150-200° C. and total pressuresin excess of 1500 kPa, the chemical reactions which occur use largequantities of oxygen, both on stoichiometric considerations and inpractice where amounts in excess of stoichiometric requirements are useddue to process inefficiencies.

An example of oxidative hydrometallurgy is the treatment of refractorygold ores or concentrates. Refractory gold ores are those gold ores fromwhich the gold cannot readily be leached by conventional cyanidationpractice. The refractory nature of these gold ores is essentially due tovery fine (sub microscopic) gold encapsulated within the sulphideminerals. This gold can often only be liberated by chemical destruction(usually oxidation) of the sulphide structure, prior to recovery of thegold, which is usually done by dissolution in cyanide solution. Ofcourse, other lixiviates such as thiourea and halogen compounds and thelike may also be used.

A number of processing options are available for treating refractorygold ores which contain sulphide minerals like pyrite, arsenopyrite andothers. Pressure oxidation, typified by the so-called Sherritt process,is one such process which typically consists of the steps of feedpreparation, pressure oxidation, solid/liquid separation, liquidneutralisation and gold recovery from oxidised solids usually bycyanidation.

A cryogenic oxygen plant is usually required to supply the substantiallevels of oxygen demand during the pressure oxidation step, which is theheart of the Sherritt process. Typically, the conditions for thepressure oxidation step require temperatures in the region of 150° C. to210° C., a total pressure of 2100 kPa, a pulp density equivalent to 20%to 30% solids by mass, and a retention time of two hours to three hours.

The typical oxidative hydrometallurgical processing methods referred toabove generally have oxidation reactions that are carried out inmulticompartment autoclaves fitted with agitators. In order to withstandthe generally highly aggressive conditions of the reactions, theautoclaves are very costly, both to install and maintain. These vesselsmust be capable of withstanding high pressure, and linings of heat andacid resistant bricks need to be used. The agitators are made oftitanium metal, and the pressure relief systems utilised are also costlyand require high maintenance. These high costs and the sophistication ofthe technology (skilled operators are generally required) detract fromthe wider acceptance of high pressure/high temperature oxidation,particularly for use in remote areas or by small to medium sizeoperators.

Cooling of the agitators also presents problems, and expensive coolingcoils and heat exchange jackets are required to keep the leachtemperature at optimum conditions.

The aggressive leaching conditions outlined for recovery of metal valuesfrom base metal concentrates are required to achieve acceptable leachingrates from the minerals. Under conditions of atmospheric pressure, theleaching rates of the mineral species are too low to support aneconomically viable leaching process.

Attempts have been made to reduce the aggressive conditions and to lowerthe pressures in order to lower the cost in building and operating aleaching plant. For instance, it is known to initially fine grind theore or the ore concentrate (it being known to use flotation as aninitial step to concentrate minerals in the ore), prior to oxidativehydrometallurgy to leach the ore. The fine grinding increases thesurface area to volume ratio of the ore particles to improve extraction.A fine grind to an 80% passing size of 15 micron or less is used. Theinitial fine grind results in acceptable leaching rates being observedwith less aggressive conditions, and leaching can be carried out attemperatures of 95-110° C. and at a pressure of about 10 atmospheres orabout 1000 kPa.

Thus, while some progress has been made in reducing the operatingparameters and thus the cost of the leach system to date, the leachstill must be carried out under pressurised conditions. Pressure leachsystems are expensive to build. Because of the high capital andprocessing costs of pressure leach systems, these systems are economicalonly for high grade concentrates. High grade concentrates are requiredbecause

(1) operating cost per unit of contained metal considerations

(2) less heat generation/exchanger problems with high grade concentrates

(3) capital cost per unit of contained metal is lower balancing up withlarge initial capacity outlay to metal recovery.

It is also known to oxidatively leach sulphide mineral species withferric ions. Ferric ion is a relatively effective oxidising agent whichenables oxidation to be carried out at pressures less than that normallyrequired when oxygen is the oxidant. However, there are a number ofpractical difficulties associated with using ferric ions as the oxidant.First, at ambient pressure the reaction is inherently slow. Also duringthe leaching reaction, ferric ions are reduced to ferrous ions. A buildup of ferrous ions in the leaching solution adversely affects the rateof leaching. Also the ferrous ions must normally be removed from theleach liquor prior to further processing which is difficult.

Leaching solutions are generally recycled. However, before a ferricleaching solution can be recycled the ferrous ions must be re-oxidisedto ferric ions. This is because it is important for maximumeffectiveness of the leach that most of the iron is in the ferric form.The leach solutions can be regenerated by electrolytic oxidation, use ofstrong oxidisers such as permanganate, oxidation under high pressure ofoxygen, or oxidation by bacteria. Each of these methods suffer fromdisadvantages which limit their application. For example high pressureoxidation is limited by the costs of the autoclaves involved. Oxidationby oxygen under ambient pressure can occur but only at an inherentlyslow rate. Catalysts may be used to increase the rate but such catalystsare expensive and are not economical for recovery from low grade ores.

Each of the above processes either require expensive autoclaves or otherequipment and/or the addition of expensive reagents used for oxidationor regeneration of ferric ions. This means that it is only economicallyviable to process high grade ores by these methods. Another disadvantageof these processes is that they generate significant amounts of wasteproducts such as gypsum, sulphuric acid and jarosite. These productsmust be disposed of in an environmentally acceptable manner which alsoadds to the cost.

Many valuable copper or zinc bearing ores are found in association withiron containing ores such as pyrite. Pyrite is of little value and iseffectively a diluent of the valuable ores. Further, leaching of pyriteproduces iron species which interfere with extraction of the desiredmetals. Pyrite is therefore generally removed from other ores prior toprocessing. The pyrite may be removed by methods such as flotation. Suchseparation also adds significantly to the cost and in some cases it isnot economically feasible to process some low grade ores at all.

OBJECT OF THE INVENTION

It is an object of the present invention to provide a method ofprocessing a mineral composition which can be carried out under mildconditions of temperature and pressure and which is economical whencompared with existing processes.

The present inventors have surprisingly discovered that by subjectingsulphide mineral compositions to fine grinding prior to leaching underconditions in which the solution chemistry is controlled in a particularmanner, such compositions can be processed under ambient conditions inopen reactors without the need for the addition of expensive reagentsand a separate step for regeneration of the leaching solution.

According to a first embodiment of the present invention there isprovided a method of processing a sulphide mineral composition which atleast partly comprises an iron containing mineral, the method comprisingthe steps of;

(a) milling said composition to a particle size P80 of 20 microns orless,

(b) leaching said composition with a solution comprising sulphuric acidand ferric ions at ambient pressure whilst sparging with an oxygencontaining gas in an open tank reactor at a temperature of up to aboutthe boiling point of the solution, whereby at least some of the acid andat least some of the ferric ions are obtained from dissolution of theiron containing mineral, and ferrous ions generated by the leachingreaction are substantially re-oxidised to ferric ions in the leachingsolution;

(c) precipitating iron and separating said iron and solid materials fromthe leaching solution;

(d) extracting desired metal ions from the leaching solution by solventextraction with an organic solvent to form an organic phase andraffinate comprising sulphuric acid and ferric ions;

(e) returning the raffinate to the open tank reactor and blending withfurther milled composition;

(f) separating the metals from the organic phase obtained in step (d) bystripping with electrolyte from an electrowinning cell andelectrowinning.

According to a second embodiment of the present invention there isprovided a method of processing a sulphide mineral composition which atleast partly comprises an iron containing mineral, the method comprisingthe steps of;

(a) milling said composition to a particle size of P80 of 20 microns orless and

(b) leaching said composition with a solution comprising sulphuric acidand ferric ions at ambient pressure whilst sparging with an oxygencontaining gas in an open tank reactor at a temperature of up to aboutthe boiling point of the solution, whereby at least some of the acid andat least some of the ferric ions are obtained from dissolution of theiron containing mineral, and ferrous ions generated by the leachingreaction are substantially re-oxidised to ferric ions in the leachingsolution.

According to a third embodiment of the present invention there isprovided a method of processing a metal sulphide flotation concentratecomprising the steps of;

(a) milling said ore to P80 of 5 micron and

(b) leaching said ore with a solution comprising sulphuric acid andferric ions, at ambient pressure whilst sparging with an oxygencontaining gas in an open tank reactor at a temperature of up to aboutthe boiling point of the solution.

The method of the present invention is applicable to any type ofsulphide mineral composition. Such compositions include ores andconcentrates. The method of the present invention is especially suitablefor processing concentrates. Examples of suitable materials includechalcopyrite, bornite, enargite, pyrite, covellite, sphalerite,chalcocite, pentlandite, cobaltite, pyrrhotite or mixtures of any two ormore thereof. Metals which can be extracted from the mineralcompositions according to the method of the first embodiment includecopper, zinc, nickel and cobalt. The concentrate grade may range fromvery low such as for example with copper containing materials 7-8 wt %copper to high grade concentrates having about 26 wt % copper.

The iron containing mineral can be any mineral which under the leachingconditions will produce ferrous or ferric ions upon dissolution.Especially preferred is pyrite, FeS₂, or pyrite ore which producesferric ion and some sulphuric acid according to the following: ##STR1##

Preferably sufficient iron containing mineral is present such that itprovides substantially all of the ferric ions in the leaching solution.The relative amounts of iron containing mineral will of course depend onthe types and amounts of the other components in the ore. Typicallyabout 20 to about 60 wt % pyrite is present. If desired, additionalpyrite or other minerals may also be added. Alternatively additionalferric sulphate may be added. It can be seen that the iron containingminerals may also provide a source of sulphuric acid. Additional acidmay need to be provided if required. Sulphuric acid is typicallygenerated in associated processes such as electrowinning and solventextraction. Preferably sulphuric acid produced in this way is recycledto the leaching step.

A preferred type of apparatus which may be suitable for producing fineor ultra fine sulphide mineral compositions in finely divided form is astirred mill. However, it will be appreciated that other types ofcomminution apparatus may also be used such as wet and dry vibratorymills or planetary mills to provide the fine or ultra fine milling ofthe invention.

In a preferred form, vertical or horizontal stirred mills generallyconsist of a tank filled with small diameter grinding media (for example1-6 mm diameter steel or ceramic balls) which are agitated by means of avertical or horizontal shaft usually fitted with perpendicular arms ordiscs. The sulphide minerals (usually contained in the form of aconcentrate) are milled by the sheering action produced by ball to ballcontact, or between balls and the stirrer or balls and the walls of thetank. The milling may be carried out dry or wet. These vertical orhorizontal stirred mills have been found to be satisfactory in providingthe required degree of fineness, and in satisfying energy and grindingmedia consumption requirements. Furthermore, the activity of the groundproduct as measured by its response to subsequent oxidation, has alsobeen found to be satisfactory. In this respect the ore is ground to amaximum average particle size of 80% passing size of 20 microns asmeasured with a laser sizer. In the present specification and claims theterm P80 is used to describe the size at which 80% of the mass of thematerial will pass. Preferably the particle size is less than P80 of 5micron. The desired particle size may vary with the type of mineralspecies used. Especially preferred paricle sizes for differentconcentrates, expressed as P80, are chalcopyrite/bornite--4.5 micron;enargite--3 micron; pyrite--3 micron; covellite--20 micron;chalcocite--20 micron; pentlandite--5 micron and cobaltite--5 micron.

The mild conditions of pressure and temperature in the oxidative leachthat follows the milling, are low when compared with the relatively highpressure and temperature conditions of known pressure oxidationtechniques such as the Sherritt process or Activox process. As indicatedabove, the Sherritt process typically requires temperatures in the orderof 150 to 210° C. and total pressures in the order of 2100 kPa. TheActivox process is designed to operate at pressures between 9 and 10atmospheres and temperatures within the range 90-110° C. However, theaccleration of the leaching response of the mineral species inaccordance with the present invention allows the oxidative leach to beconducted at temperatures below about 100° C. and at atmosphericpressure in cheap open tank reactors.

With the preferred operating conditions being at about 60° C. up to theboiling point of the solution and at 1 atm total pressure, a low costreactor such as an open tank is sufficient to serve as the leachingvessel. There also is no need for the use of expensive titanium metalagitators due to the less corrosive nature of the leach solution.Furthermore, abrasion problems are substantially reduced due primarilyto the fine nature of the feed.

Importantly, the complex heat exchange and pressure let down systemsnecessary for operation of a pressure vessel are not necessary as thereactor operates at atmospheric pressure. Excess heat is removed fromthe system through solution evaporation and this removes the need forcostly heat exchange facilities. Also the reaction becomes autogenous atabove about 60 to about 70° C. If additional heating is required thiscan be easily done by known methods such as the injection of steam.

Further, much higher percent solids slurries can be treated by thedescribed method due to the relaxation of the requirements for lowsulphur levels in feed to an autoclave necessary for heat controlpurposes. Typically the leach slurry density varies from about 10 toabout 20 wt %.

In the method of the first embodiment the leaching solution is typicallyrecycled raffinate from the solvent extraction step. In this casepreferably most of the ferric ion and sulphuric acid may be generated bythe leach/solvent extraction/electrowinning process. If the leach is notpart of a continuous process, ferric and sulphuric acid may be added ifrequired. Typically the raffinate comprises 30-40 g/l H₂ SO₄ and 10-20g/l Fe. The Fe will normally be present as a mixture of ferric andferrous ions.

The leaching solution is sparged with an oxygen containing gas. The gasmay be air or preferably oxygen or oxygen enriched air. The gas flow isdependent upon the amount of oxygen required to sustain the leachingreaction and regeneration of the ferric ions. Typically the gas flow isabout 400 to about 1000 kg O₂ per ton of metal produced. If desired asurfactant or the like may be added to minimize frothing of the leachingsolution. Under the conditions of the leaching reaction, the metals maybe oxidised by ferric ion according to the following general reaction:

    MS+2Fe.sup.3+ →M.sup.2+ +2Fe.sup.2+ +S°

Further oxidation of the elemental sulphur to sulphate according to thereaction:

    S°+3/2O.sub.2 +H.sub.2 O→H.sub.2 SO.sub.4

requires elevated temperature and pressure and does not occur to anysignificant extent under the leaching conditions of the presentinvention. For example, at 90° C. at atmospheric pressure, in theabsence of bacterial catalysis, less than 5% of the elemental sulphur isoxidised to sulphate. By comparison, at 180° C. and 12 atmospheresoxygen partial pressure, most of the sulphur is oxidised to sulphate.Oxidation to sulphate has several disadvantages as additionalneutralising reagents are required during the postleachingneutralisation steps. A further advantage of formation of elementalsulphur is that gaseous emission such as sulphur dioxide is minimizedwhich causes an environmental hazard. Further by not carrying out theoxidation completely to sulphate, the consumption of oxygen issignificantly reduced which saves on operating costs. For exampleconventional PSA oxygen plants may be sufficient to supply the oxygenwithout the need for cryogenic oxygen plants. This in turn reduces thecapital cost and the operating costs by using equipment that is simpleto operate.

The ferric ions are regenerated by reaction of the ferrous ions withoxygen according to:

    4Fe.sup.2+ +4H.sup.+ +O.sub.2 →4Fe.sup.3+ +2H.sub.2 O

Typically oxidation of the ferrous ion occurs at a rate of 2-5 g offerrous ion oxidised per litre of slurry per hour of reaction.

After substantially all the mineral has been oxidised the leach slurrymay be further processed according to known methods. Preferably theslurry is filtered to remove solids and the clear liquid subjected tosolvent extraction followed by electrowinning. Typically the leachslurry is neutralised prior to any further processing. As describedabove the production of sulphate is reduced under the conditions of thepresent invention, thereby minimising the amount of neutralising agentswhich need to be added. Typically the slurry is neutralised by theaddition of limestone or the like. This also precipitates excess iron,arsenic and other impurities generated in the leach. Tests carried outunder the conditions of the present invention have also indicated thatiron can be selectively precipitated and remains in the leach residue asgoethite, jarosite or some form of hydrated oxide, whilst valuableminerals like nickel, copper or zinc remain in solution. If desired theprecipitated solids may be further filtered and any remaining liquid maybe returned to the leach solution for further processing. As describedabove substantially all of the sulphide sulphur is oxidised to elementalsulphur during the leaching reaction. The elemental sulphur is presentas finely dispersed granules. Because the leach is conducted attemperatures below the melting point of sulphur ie 120° C.,agglomeration of molten sulphur is avoided. The granulated sulphur isnormally removed from the leaching solution with the goethite and/orother iron residue.

The solids which are separated from the leaching solution may be furthertreated to extract any precious metals such as gold, platinum or silver.These methods of extraction such as cyanidation for gold are well knownin the art.

The steps of solvent extraction and electrowinning are well known in theart and need not be described in detail. Typically the neutralisedslurry may filtered and extracted with an organic solvent which recoversmetals such as copper, nickel or zinc. The metals may then be strippedfrom the organic phase by known methods. The metals are then separatedfrom the electrolyte by electrowinning. The spent electrolyte may thenbe returned to the stripping stage.

BRIEF DESCRIPTION OF THE DRAWINGS

FIG. 1 is a flow diagram of the method according to a preferredembodiment of the invention.

BEST MODE

The present invention will now be described in relation to the followingexamples. However, it will be appreciated that the generality of theinvention as described above is not to be limited by the followingdescription.

EXAMPLE ONE Enargite and Chalcocite Leaching

Copper flotation concentrate containing 19.5% copper, 4.0% arsenic, 23%iron, 2.35 g/t gold and 36% sulfide sulfur was milled to a size of 80%passing 5 micron in a horizontal 1 litre stirred ball mill.Mineralogically the concentrate was composed of 11.9% chalcocite (Cu₂S), 20.9% enargite (Cu₃ AsS₄), 50% pyrite (FeS₂) and the remainder wassiliceous gangue minerals.

The milled pulp was leached at 90° C. in an open reactor utilising aferric ion sulfuric acid lixiviant solution with either oxygen or airsparging. The solids were recovered by filtration and reslurry washedwith 5% v/v sulfuric acid solution, prior to being dried and assayed.Post leach solutions were analysed for copper, arsenic and iron byconventional atomic absorption spectroscopy analytical methods. Ferrousand ferric levels were determined by potassium permanganate titrations,whilst the acid levels were determined by a neutralisation method.

Greater than 92% copper dissolution was achieved from concentrate milledto a particle size of P₈₀ 3.5 microns in 10 hours employing leachconditions of 10% pulp density, 30 g/L ferric ions, 50 g/L sulfuricacid, 90° C., oxygen sparging and 2.0 kg/ton lignosol. Lignosol was usedto reduce the amount of frothing in the initial stages of the leach.

The leach residue was then leached in aerated sodium cyanide solution torecover gold. The leach test was conducted at ambient temperature with300 ppm free NaCN maintained in the leach solution during the leach. Thetest was fun for 24 hours at a pulp density of 45% solids. Greater than82% gold extraction was achieved from the residue.

The predominate leaching mechanism thought to be occurring in the abovesystem involves ferric ions acting as the oxidant, although it isconceivable that an acid/oxygen mechanism is also operating. Thepredominate reactions occurring in this leaching system are presentedbelow. ##STR2##

Ferric iron was regenerated in the leach solution by the action ofoxygen on ferrous iron according to the reaction:

    2FeSO.sub.4 +H.sub.2 SO.sub.4 +1/20.sub.2 →Fe.sub.2 (SO).sub.3 +H.sub.2 O

In this way, the ferric oxidant was continually regenerated during theleaching process.

The pulp density in the reactor appeared to be limited by the solubilityof copper sulfate in the resulting iron/acid electrolyte. The use of airinstead of oxygen increased the leaching residence time from 10 hours(oxygen) to 14 hours (air) without loss of overall copper recovery.

EXAMPLE TWO Chalcocite Leaching

Copper flotation concentrate containing 8.1% copper, 0.2% arsenic, 13.8%iron and 18% sulfide sulfur was milled to a size of 80% passing 5 micronin a horizontal 1 litre stirred ball mill. This concentrate contained9.4% chalcocite (Cu₂ S), 1.3% enargite (Cu₃ AsS₄), 29.6% pyrite (FeS₂)and the remainder was siliceous gangue.

The milled pulp was leached at 90° C. in an open reactor utilising aferric ion sulfuric acid lixiviant solution with either oxygen or airsparging. Greater than 95% copper dissolution was achieved in 10 hoursemploying leach conditions of 10% pulp density, 30 g/L ferric ions, 50g/L sulfuric acid, 90° C., oxygen sparging and 2.0 kg/ton lignosol.

EXAMPLE THREE Chalcopyrite Leaching

Copper flotation concentrate containing 18.0% copper, 25.5% iron and18.6% sulfide sulfur was milled to a size of 80% passing 5 micron in ahorizontal 1 litre stirred ball mill. This concentrate contained 51.8%chalcopyrite (CuFeS₂), 20.8% pyrite (FeS₂) and the remainder wassiliceous gangue.

The milled pulp was leached at 80° C. in an open reactor utilising aferric ion sulfuric acid lixiviant solution with oxygen sparging.Greater than 95% copper dissolution was achieved in 10 hours employingleach conditions of 10% pulp density, 5 g/L ferric ions, 20 g/l ferrousiron, 50 g/L sulfuric acid, 90° C., oxygen sparging and 2.0 kg/tonlignosol.

The method of the invention can be used in association with otherupstream or downstream processes. For instance, prior to the fine grindand leach, the ore can be treated in one or more flotation steps.Downstream processes can include flotation processes with or withoutadditional fine grinding, and/or can include solvent extraction andelectrowinning steps.

EXAMPLE FOUR Nickel Leaching

Nickel bearing concentrate containing 1.7% nickel, 0.03% cobalt, 11%iron and 16% sulfide sulfur was milled to a size of 80% passing 5 micronin a horizontal 1 litre stirred ball mill. Mineralogically theconcentrate was composed of pentlandite, pyrite and the remainder wassiliceous gangue materials.

The milled pulp was leached at 90° C. in an pen reactor utilising aferric ion sulfuric acid lixiviant solution with oxygen sparging. Thesolids were recovered by filtration and reslurry washed with water,prior to being dried and assayed. Post leach solutions were analysed fornickel, cobalt and iron by conventional atomic absorption spectroscopyanalytical methods. Ferrous and ferric levels were determined bypotassium permanganate titrations, whilst the acid levels weredetermined by a neutralisation method.

Greater than 92% nickel and 86% cobalt dissolution was achieved fromconcentrate milled to a particle size P80 of 5 microns in a leach timeof 8 hours employing leach conditions of 10% pulp density, 5 g/L ferricions, 20 g/l ferrous iron, 80 g/L sulfuric acid, 90° C., oxygen spargingand 2.0 kg/ton lignosol. Lignosol was used to reduce the amount offrothing in the initial stages of the leach.

EXAMPLE FIVE Cobalt Leaching

Cobalt bearing concentrate containing 0.309% cobalt, 8.5% iron and 0.66%arsenic was milled to a size of 80% passing 3 microns in a horizontal 1litre stirred ball mill. Mineralogically the concentrate was composed ofcobaltite and cobaltiferous pyrite, pyrite and the remainder wassiliceous gangue minerals.

The milled pump was leached at 90° C. in an open reactor utilising aferric ion sulfuric acid lixiviant solution with oxygen sparing. Thesolids were recovered by filtration and reslurry washed with water,prior to being dried and assayed. Post leach solutions were analysed forcobalt and iron by conventional atomic absorption spectroscopyanalytical methods. Ferrous and ferric levels were determined bypotassium permanganate titrations, whilst the acid levels weredetermined by a neutralisation method.

Greater than 79% cobalt dissolution was achieved from concentrate milledto a particle size of P80 of 3 microns in a leach time of 8 hoursemploying leach conditions of 10% pulp density, 10 g/L ferric ions, 50g/L sulfuric acid, 90° C., oxygen sparging and 2.0 kg/ton lignosol.

EXAMPLE SIX Zinc Leaching

Zinc concentrate containing 46.6% zinc, 10% iron and 2.8% lead wasmilled to a size of 80% passing 3 micron in a horizontal 1 litre stirredball mill. Mineralogically the concentrate was composed of sphalerite,galena, pyrite and the remainder was siliceous gangue minerals.

The milled pulp was leached at 90° C. in an open reactor utilising aferric ion sulfuric acid lixiviant solution with oxygen sparging. Thesolids were recovered by filtration and reslurry washed with water,prior to being dried and assayed. Post leach solutions were analysed forzinc and iron by conventional atomic absorption spectroscopy analyticalmethods. Ferrous and ferric levels were determined by potassiumpermanganate titrations, whilst the acid levels were determined by aneutralisation method.

Greater than 97% zinc extraction was achieved from zinc concentratemilled to a particle size of P80 of 3 microns in a leach time of 8 hoursemploying leach conditions of 10% pulp density, 10 g/L ferric ions, 50g/L sulfuric acid, 90° C., oxygen sparging and 2.0 kg/ton lignosol.

EXAMPLE SEVEN Chalcopyrite Leaching as a Continuous Process

The example below describes the operation of a fully continuous pilotplant designed to produce 8 kg per day of LME grade A cathode copperfrom a copper concentrate. The pilot plant ran for 21 days treating thefeed outlined below.

Referring to FIG. 1 a concentrate sample of the composition listed belowwas slurried in tap water at a slurry density of 60% w/w. The slurry wasthen milled in step 1 to a particle size of 80% passing 10 microns in ahorizontal stirred bead mill.

Table 1 Composition of the concentrate sample:

    ______________________________________                                               CuFeS.sub.2                                                                         37% w/w                                                            FeS.sub.2 44% w/w                                                             SiO.sub.2 11% W/W                                                             Other  8% W/W                                                               ______________________________________                                    

The slurry sample was then mixed in step 2 with solvent extraction plantraffinate 3 to dilute the slurry density of 15% w/w. The raffinatecontained 35 g/l H₂ SO₄, 9 g/l ferric iron and 10 g/l ferrous iron.

The diluted slurry was then pumped through a three vessel leaching train4,5,6 at a flow rate designed to give a residence time in the leachingtrain of 20 hours. The leaching train consisted of three agitated 100litre baffled tanks. Slurry flowed by gravity from one tank to the next.The tanks were maintained at 90° C. by a combination of the exothermicnature of the leach reaction and the injection of live steam into theslurry. Oxygen was injected into the slurry by air spear 7,8,9 locatedbelow the leach agitator. The oxygen was added at a rate of 600 kg perton of copper produced. The copper extraction across the leachingcircuit was 97% w/w.

No acid or iron sulphate was added to the leaching circuit for theduration of the 21 day pilot plant run.

Leached slurry overflowed the final leach tank into a neutralisationtank 10 and then to a thickener 10A. The leach solution composition wastypically 17-19 g/l copper and 35-45 g/l iron. The slurry wasneutralised to pH 2.0 with limestone slurry 12 to precipitate iron fromthe leach slurry as goethite 13. Finely granulated elemental sulphur wasremoved with the goethite. The slurry comprising the goethite, sulphurand leach residue was filtered 11 and any liquid was returned to theleach solution 14.

Neutralised slurry was then pumped through a plate and frame pressurefilter 15. The filtrate contained 17-19 g/l copper, 20 g/l iron and 5g/l H₂ SO₄. The filtrate was then pumped through a three stage solventextraction plant 16,17,18 to recover copper from the leach liquor. Theraffinate from the solvent extraction stage contained 0.3 g/l copper, 20g/l iron and 35 g/l H₂ SO₄ and 20 g/l iron, and was transferred back tothe leaching circuit and blended with more ground concentrate at 2.

Loaded organic from the solvent extraction plant was then stripped 19with spent electrolyte containing 180 g/l H₂ SO₄. The stripped organic20 was pumped back to the extraction stage.

The rich electrolyte 21 was then pumped through a 200 litreelectrowinning cell 22 containing two cathodes and three anodes. Eachcathode face had an area of 0.25 m². Copper was plated out from the richelectrolyte at a current density of 280 A/m² to produce a cathode plate.The spent electrolyte from the cell contained 180 g/l H₂ SO₄ and 32 g/lcopper. The copper cathode was analysed and met the requirements for LMEGrade A.

The goethite leach residue was leached in sodium cyanide to determinethe amount of gold which could be recovered from the oxidisedconcentrate. The leach residue was leached at 45% w/w solids at pH 10,with a free cyanide level of 300 ppm maintained throughout the leach.The slurry was leached for 24 hours in a batch test. Gold recovery fromthe leach residue was 92.5% w/w.

It can be seen that the methods of the present invention offer a numberof advantages over existing methods. The need for tight control onconcentrate grade is relaxed due to the lower operating costs of theleach relative to a pressurized leach, and the control of excess heatgeneration through evaporative cooling of the open tank reactors.

The leach can be conducted in cheap open tanks instead of expensivepressure vessels. An equivalent sized pressure leach facility may costabout 6 to 8 times as much as the open tank leaching system of thepresent invention. This also enables a leaching circuit to beconstructed practically on site. This avoids costs of transportationwhich can be considerable. In some cases these costs may make ituneconomical to transport and process low grade ores.

The leach circuit is less sensitive to grade of metal. Therefore thiswill enable higher metal recovery in upstream unit operations.

The method of the present invention is also capable of producing a highgrade electrowin metal. For example in some cases it may be possible toproduce a product through solvent extraction/electrowinning which can besold directly. Smelted metal generally requires further refining. Thisenables substantial cost savings as well as producing a product whichwill attract a premium price.

The leach reaction is self-sustaining when coupled with a solventextraction-electrowinning plant, as the only reagents that need to beadded to the leach are air/oxygen and a neutralising agent such as lime.The need for addition of expensive reagents is eliminated.

The present invention is ideal for mineralogically complex ores whichare finely disseminated base metal sulphides minerals with other sulfide(e.g. chalcocite, sphalerite, enargite covering pyrite ores) that havebeen traditionally difficult to treat metallurgically.

It should be appreciated that various other changes and modificationsmay be made to the embodiments described without departing from thespirit and scope of the invention.

We claim:
 1. A method of processing a sulphide mineral composition whichat least partly comprises an iron containing mineral, the methodcomprising the steps of;(a) milling said composition to a particle sizeP80 of 20 microns or less, (b) leaching said composition with a solutioncomprising sulphuric acid and ferric ions at ambient pressure whilstsparging with an oxygen containing gas in an open tank reactor at atemperature of up to about the boiling point of the solution, wherein atleast some of the acid and at least some of the ferric ions are obtainedfrom dissolution of the iron containing mineral, and ferrous ionsgenerated by the leaching reaction are substantially re-oxidised toferric ions in the leaching solution; (c) precipitating iron andseparating said iron and solid materials from the leaching solution; (d)extracting desired metal ions from the leaching solution by solventextraction with an organic solvent to form an organic phase andraffinate comprising sulphuric acid and ferric ions; (e) returning theraffinate to the open leaching tank reactor and blending with furthermilled composition; (f) separating the metals from the organic phaseobtained in step (d) by stripping with electrolyte from anelectrowinning cell and electrowinning.
 2. The method of claim 1,wherein said composition is a flotation concentrate.
 3. The method ofclaim 2, wherein said concentrate is a low grade concentrate.
 4. Themethod of claim 1, wherein said iron containing mineral is a pyrite ore.5. The method of claim 4, wherein said mineral composition comprises 20to 60 wt % pyrite.
 6. The method of claim 5, wherein substantially allof said ferric ions are generated by dissolution of pyrite.
 7. Themethod of claim 1, wherein said composition is milled to a particle sizeof P80 of 10 micron or less.
 8. The method of claim 1, wherein saidtemperature is from about 60° C. up to about the boiling point of theleaching solution.
 9. The method of claim 1, wherein said gas is oxygen.10. The method of claim 9, wherein oxygen is sparged at a rate of 400 to1000 kg per ton of metal produced.
 11. The method of claim 1, whereinthe solid from step (c) is further leached to recover any preciousmetals.
 12. The method of claim 11, wherein said precious metals areselected from the group consisting of gold, platinum or silver.
 13. Themethod of claim 1, wherein said metal is selected from the groupconsisting of copper, zinc, nickel or cobalt.
 14. A method of processinga sulphide mineral composition comprising about 30 to about 40 wt %chalcopyrite, about 40 to about 50 wt % pyrite and up to 20 wt %siliceous gangue, the method comprising(a) milling the composition to aparticle size P80 of 10 microns, (b) leaching said composition with asolution comprising sulphuric acid and ferric ions at ambient pressurewhilst sparging with oxygen at a rate of about 600 kg per ton of copperproduced in an open reactor at a temperature of about 90° C., (c)neutralizing the leaching solution with limestone to precipitate excessiron as goethite and removing the goethite and any other solids from theleaching solution; (d) filtering the leaching solution and extractingdissolved copper from the leaching solution by solvent extraction withan organic solvent, such that the raffinate comprises sulphuric acid andferric ions; (e) returning the raffinate to the leaching tank andblending with further milled composition; (f) separating the copper fromthe organic phase obtained in step (c) by stripping with electrolytefrom an electrowinning cell and electrowinning.
 15. A method ofprocessing a sulphide mineral composition which at least partlycomprises an iron containing mineral, the method comprising the stepsof;(a) milling said composition to a particle size P80 of 20 microns orless and (b) leaching said composition with a solution comprisingsulphuric acid and ferric ions at ambient pressure whilst sparging withan oxygen containing gas in an open tank reactor at a temperature of upto about the boiling point of the solution, whereby at least some of theacid and at least some of the ferric ions are obtained from dissolutionof pyrite, and ferrous ions generated by the leaching reaction aresubstantially re-oxidised to ferric ions in the leaching solution. 16.The method of claim 15 wherein said composition is a flotationconcentrate.
 17. The method of claim 15 wherein said iron containingmineral is a pyrite ore.
 18. The method of claim 17 wherein saidcomposition comprises 20 to 60 wt % pyrite.
 19. The method of claim 18wherein substantially all of said ferric ions are generated bydissolution of pyrite.
 20. The method of claim 15 wherein saidcomposition is milled to a particle size P80 of 5 micron.
 21. The methodof claim 15 wherein said temperature is about 60° C. up to the boilingpoint of the leaching solution.
 22. The method of claim 15 wherein saidgas is oxygen.
 23. The method of claim 22 wherein oxygen is sparged at arate of about 400 to about 1000 kg per ton of metal produced.
 24. Themethod of claim 15 wherein a surfactant is added in step (a) to minimizefrothing of the solution.
 25. A method of processing a metal sulphideflotation concentrate comprising the steps of;(a) milling said ore toP80 of 5 micron and (b) leaching said concentrate with a solutioncomprising sulphuric acid and ferric ions, at ambient pressure whilstsparging with an oxygen containing gas in an open tank reactor at atemperature of up to about the boiling point of the solution.
 26. Themethod of claim 25 wherein said temperature is about 60° C. up to theboiling point of the leaching solution.
 27. The method of claim 26wherein said gas is oxygen.
 28. The method of claim 27 wherein oxygen issparged at a rate of about 400 to about 1000 kg per ton of metalproduced.
 29. The method of claim 28 wherein a surfactant is added instep (a) to minimise frothing of the solution.